Process for Leaching Metal Sulfides with Reagents Having Thiocarbonyl Functional Groups

ABSTRACT

This application pertains to methods of recovering metals from metal sulfides that involve contacting the metal sulfide with an acidic sulfate solution containing ferric sulfate and a reagent that has a thiocarbonyl functional group, wherein the concentration of reagent in the acidic sulfate solution is sufficient to increase the rate of metal ion extraction relative to an acidic sulfate solution that does not contain the reagent, to produce a pregnant solution containing the metal ions.

This application claims priority to U.S. Patent Application No.62/149,015, filed Apr. 17, 2015, the contents of which are incorporatedherein by reference.

BACKGROUND 1. Field of the Disclosure

This disclosure pertains to methods for leaching metals from metalsulfide-containing ores. More particularly it pertains to ahydrometallurgical process for the extraction of metals from metalsulfide-containing ores using reagents having a thiocarbonyl functionalgroup.

2. Description of Related Art

Aqueous processing of minerals presents several advantages overpyrometallurgical approaches, particularly when dealing with complexand/or low-grade ores. The main disadvantage of hydrometallurgicalprocesses, when applied to several metal sulfide ores, is the lowextraction rates that are observed. It is desirable to develop a processwhere high metal extractions can be achieved in time scales that are ofindustrial interest.

Chalcopyrite, for example, is a semiconductor, and therefore corrodeselectrochemically in oxidizing solutions. In ferric sulfate media, theoverall leaching reaction is as follows:

CuFeS₂(s)+2Fe₂(SO₄)₃(a)→CuSO₄(a)+5FeSO₄(a)+2S⁰(s)

This reaction may be represented as a combination of anodic and cathodichalf-cell reactions:

Anodic half-cell reaction: CuFeS₂→Cu²⁺+Fe²⁺+2S⁰+4e ⁻

Cathodic half-cell reaction: 4Fe³⁺+4e ⁻→4Fe²⁺

A fundamental problem with chalcopyrite oxidation is that chalcopyritemineral surfaces become resistant to electrochemical breakdown atsolution potentials above a certain level (generally considered to beabout 550 to 600 mV vs Ag/AgCI). It is widely held that this resultsfrom the formation of some sort of passivating film on the mineralsurface that most likely consists of an altered, partially Fe-depletedform of chalcopyrite. It is desirable to provide leaching processes inwhich such passivation is reduced or avoided.

Some work has been done in extractive hydrometallurgy to recoverprecious metals such as gold and silver from copper concentrates orchalcopyrite residues after copper extraction. Deschênes and Ghali(Hydrometallurgy 20:129-202) demonstrated the potential application ofthiourea in acidic sulfate leaching of sulfide concentrates, such asthose containing chalcopyrite, to selectively recover gold and silver.Thiourea is an organosulfur compound having a thiocarbonyl functionalgroup. However, thiourea did not appear to have an effect on therecovery of copper from copper sulfides.

SUMMARY

This disclosure relates, at least in part, to the unexpected discoverythat several reagents comprising a thiocarbonyl function group (e.g.thiourea) can be used to facilitate the leaching of metal from severalmetal sulfides (e.g. copper from chalcopyrite) with acidic sulfate leachsolutions. When added in small amounts, such reagents may increase therate of metal leaching over that observed in its absence.

This disclosure relates to a method of recovering at least one metalfrom an ore containing at least one metal sulfide. The method involvescontacting the ore with an acidic sulfate solution containing ferricsulfate and a reagent having a thiocarbonyl functional group to extractmetal ions from the at least one metal sulfide, wherein theconcentration of the reagent in the acidic sulfate solution issufficient to increase the rate of the metal ion extraction relative toan acidic sulfate solution that does not contain the reagent, to producea pregnant solution containing the metal ions. The method furtherinvolves recovering the at least one metal from the pregnant solution.The at least one metal includes: copper, wherein the at least one metalsulfide includes chalcopyrite, covellite, bornite, enargite, a coppersulfide of the formula Cu_(x)S_(y) wherein the x:y ratio is between 1and 2, or a combination thereof; cadmium, wherein the at least one metalsulfide is greenockite; nickel, wherein the at least one metal sulfideis pentlandite, violarite, or a combination thereof; or a combinationthereof. The concentration of the reagent may be in the range of about0.2 mM to about 30 mM.

This disclosure also relates to a method of recovering at least onemetal from an ore containing at least one metal sulfide. The methodinvolves contacting the metal sulfide with an acidic sulfate solutioncontaining a reagent having a thiocarbonyl functional group, wherein theinitial concentration of the reagent in the acidic sulfate solution isin the range of about 0.2 mM to about 30 mM or less, to produce apregnant solution containing metal ions. The method further involvesrecovering copper from the pregnant solution. The at least one metalincludes: copper, wherein the at least one metal sulfide includeschalcopyrite, covellite, bornite, enargite, a copper sulfide of theformula Cu_(x)S_(y) wherein the x:y ratio is between 1 and 2, or acombination thereof; cadmium, wherein the at least one metal sulfide isgreenockite; nickel, wherein the at least one metal sulfide ispentlandite, violarite, or a combination thereof; or a combinationthereof.

In the methods described above, the concentration of the reagent in theacidic sulfate solution may be in the range of about 0.2 mM to about 20mM, about 0.2 mM to about 10 mM, about 0.2 mM to about 5 mM, about 0.2mM to about 4 mM, about 0.2 mM to about 3 mM, about 0.2 mM to about 2mM, about 0.2 mM to about 1.5 mM, about 0.2 mM to about 1.0 mM, or about0.2 mM to about 0.5 mM.

Where the metal is a copper sulfide of the formula CuxSy wherein the x:yratio is between 1 and 2, the copper sulfide may includes chalcocite,djurleite, digenite, or a combination thereof.

In the methods described above, the reagent may be thiourea (Tu),thioacetamide (TA), sodium-dimethyldithiocarbamate (SDDC), ethylenetrithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combinationthereof.

This disclosure yet further relates to a method of recovering a metalfrom an ore containing at least one metal sulfide. The method involvescontacting the ore with an acidic sulfate solution comprising ferricsulfate and formamidine disulfide (FDS) to produce a pregnant solutioncontaining metal ions. The method further involves recovering the metalfrom the pregnant solution. The at least one metal includes: copper,wherein the at least one metal sulfide includes chalcopyrite, covellite,bornite, enargite, a copper sulfide of the formula Cu_(x)S_(y) whereinthe x:y ratio is between 1 and 2, or a combination thereof; cadmium,wherein the at least one metal sulfide is greenockite; nickel, whereinthe at least one metal sulfide is pentlandite, violarite, or acombination thereof; or a combination thereof. The concentration of FDSin the acidic sulfate solution may be in the range of about 0.1 mM toabout 15 mM, about 0.1 mM to about 10 mM, about 0.2 mM to about 5 mM,about 0.1 mM to about 2.5 mM, about 0.1 mM to about 2 mM, about 0.1 mMto about 1.5 mM, about 0.1 mM to about 1.0 mM, about 0.1 mM to about 0.5mM, or about 0.1 mM to about 0.25 mM. Where the metal is a coppersulfide of the formula CuxSy wherein the x:y ratio is between 1 and 2,the copper sulfide may includes chalcocite, djurleite, digenite, or acombination thereof.

The concentration of FDS in the acidic sulfate solution may besufficient to provide sufficient thiourea to increase the rate of themetal ion extraction relative to an acidic sulfate solution that doesnot contain the reagent to produce the pregnant solution containing themetal ions.

In the methods described above, wherein the ore may be provided ascoarse particles, which may be agglomerated particles. Ferric ions maybe used to oxidize the metal sulfide. In the methods described above,the ferric ions may be generated at least in part by bacteria.

The methods may involve a percolation leach. The percolation leach maybe a heap leach. The percolation leach may be a vat leach. The leach maybe a tank leach.

Recovering metal from the pregnant solution may include solventextraction and electrowinning.

Other aspects and features of the present invention will become apparentto those ordinarily skilled in the art upon review of the followingdescription of specific embodiments of the invention in conjunction withthe accompanying figures.

BRIEF DESCRIPTION OF THE DRAWINGS

In drawings which illustrate embodiments of the invention,

FIG. 1 is a plot showing the effect of thiourea concentration on mixedpotential and dissolution current density (i_(dissol)) of the CuFeS₂electrode;

FIG. 2 is a bar graph showing electrochemical dissolution rates of aCuFeS₂ electrode in sulfuric acid solution at pH 2 and 25° C. withvarying initial concentrations of thiourea, formamidine disulfide (FDS),and Fe(III);

FIG. 3 is a schematic diagram for the leaching column used in respect ofthe leaching experiments pertaining to FIGS. 4, 5, and 6;

FIG. 4 is a graph showing the effect of thiourea concentration on theleaching of copper from Ore A in column leach experiments;

FIG. 5 is a graph showing the effect of thiourea concentration on theleaching of copper from Ore B in column leach experiments;

FIG. 6 is a graph showing the effect of thiourea concentration on theleaching of copper from Ore C in column leach experiments;

FIG. 7 is a graph showing the effect of thiourea concentration on theleaching rate of copper from Ore C in column leach experiments;

FIG. 8 is a graph showing the effect of thiourea concentration on ORPover time;

FIG. 9 is a graph showing the effect of thiourea concentration on copperdissolution for coarse Ore A in bottle roll experiments;

FIG. 10 is a graph showing the effect of thiourea concentration oncopper dissolution for coarse Ore B in bottle roll experiments;

FIG. 11 is a graph showing the effect of Tu addition on various mineralsthat contain Cu(I). Diamonds pertain to bornite, triangles refer tocovellite, inverted triangles pertain to chalcocite, and squares pertainto chalcopyrite. Open symbols refer to control treatments without Tu,whereas solid symbols refer to minerals treated solutions having aninitial Tu concentration of 2 mM;

FIG. 12 is a graph showing the effect of Tu on cadium extraction fromgreenockite;

FIG. 13 is a graph showing the effect of Tu on copper extraction fromenargite;

FIG. 14 is a graph showing the effect of Tu on nickel extraction fromviolarite;

FIG. 15 is a graph showing the percentage of Cu ions remaining insolution after various amounts of Tu addition;

FIG. 16 is a graph showing extraction of Cu from chalcopyrite undervarious Tu dosages;

FIG. 17 is a graph showing the relationship between Tu dosage and Cuextraction after 172 hours;

FIG. 18 is a graph showing leaching of copper from chalcopyrite instirred reactor tests using reagents comprising thiocarbonyl functionalgroups. Circles pertain to Tu, triangles pertain to TA, invertedtriangles pertain to SDDC, diamonds pertain to ETC, stars pertain toTSCA, and squares pertain to controls;

FIG. 19 is a graph showing leaching of copper from covellite in stirredreactor tests using reagents comprising thiocarbonyl functional groups.Circles pertain to Tu, triangles pertain to TA, diamonds pertain toSDDC, and squares pertain to controls;

FIG. 20 is a graph showing leaching of copper from bornite in stirredreactor tests using reagents comprising thiocarbonyl functional groups.Triangles pertain to Tu, circles pertain to TA, and squares pertain tocontrols;

FIG. 21 is a graph showing leaching of copper from enargite in stirredreactor tests using reagents comprising thiocarbonyl functional groups.Circles pertain to Tu, triangles pertain to TA, inverted trianglespertain to ETC, and squares pertain to controls;

FIG. 22 is a graph showing the leaching of copper from chalcopyrite instirred reactor tests using reagents comprising thiocarbonyl functionalgroups, urea, and carbon disulfide. Circles pertain to urea, trianglespertain to controls, inverted triangles pertain to TA, diamonds pertainto Tu, stars pertain to ETC, and squares pertain to carbon disulfide;

FIG. 23, panel a, is a graph comparing the leaching of copper fromchalcopyrite (circles) or bornite (triangles) using leaching solutionswith either an initial concentration of 2 mM Tu (solid symbols) or aninitial concentration of 1 m FDS (open symbols);

FIG. 23, panel b, is a graph comparing the leaching of copper fromcovellite (circles) or chalcocite (triangles) using leaching solutionswith either an initial concentration of 2 mM Tu (solid symbols) or aninitial concentration of 1 m FDS (open symbols);

FIG. 24 is a graph monitoring bacterial activity and FDS content withORP and HPLC; and

FIG. 25 is a graph showing the bioleaching of CuFeS₂ using only Fe³⁺(day 0-50) and using Fe³⁺+Tu (day 90-150) in closed loop experiments.

DETAILED DESCRIPTION

This disclosure relates to methods of recovering metal from a metalsulfide mineral, and relates in particular to the unexpected discoverythat various reagents having a thiocarbonyl functional group, e.g.thiourea (also known as thiocarbamide), can be used to facilitate theleaching of metal from a metal sulfide mineral with acidic sulfate leachsolutions. Such reagents can increase the rate of metal sulfideleaching.

Such methods may be particularly useful in the recovery of metal fromlow grade ores that do not contain the metal sulfide mineral in highproportions. The method involves contacting the copper sulfide mineralwith an acidic sulfate solution containing the reagent having athiocarbonyl functional group.

Minerals Chalcopyrite (CuFeS₂)

The leaching of chalcopyrite is accomplished in acidic ferric sulfatesolution according to the following reaction formula:

CuFeS₂+4Fe³⁺→Cu²⁺+5Fe²⁺+2S⁰

Covellite (CuS)

Leaching of covellite in ferric sulfate solution proceeds according tothe following reaction formula:

CuS+2Fe³⁺→Cu²⁺+2Fe²⁺+S⁰

Chalcocite (Cu₂S)

Leaching of chalcocite in ferric solution proceeds according to thefollowing formula:

Cu₂S+2Fe³⁺→Cu²⁺+2Fe²⁺+CuS

The skilled person understands that that “chalcocite” ores frequentlycontain a mixture of minerals with the formula Cu_(x)S_(y), where thex:y ratio is between 1 and 2. Additional minerals within this formulainclude digenite and djurleite.

Bornite (Cu₅FeS₄)

Bornite is an important copper mineral that usually coexists withchalcopyrite. The leaching process of bornite in ferric solution isdescribed in two stages:

CusFeS₄+4Fe³⁺→Cu₃FeS₄+2Cu²⁺+4Fe²⁺

Cu₃FeS₄+8Fe³⁺→3Cu²⁺+9Fe²⁺+4S⁰

Enargite (Cu₃AsS₄)

Unlike the other copper minerals mentioned above (chalcopyrite,covellite, charcocite and bornite), the copper in enargite is mainlyCu(II) instead of Cu(I). The difference in copper's oxidation state willalso influence its leaching kinetics under catalyzed conditions.Previous study showed that the leaching of enargite at atmosphericpressure is extremely slow. The dissolution of enargite in ferricsulfate media can take various paths. Two of them are described asfollows:

Cu₃AsS₄+20H₂O+35Fe³⁺→3Cu²⁺+AsO₄ ³⁻+4SO₄ ²⁻+40H⁺+35Fe²⁺

Cu₃AsS₄+4H₂O+11Fe³⁺→3Cu²⁺+AsO₄ ³⁻+4S⁰+8H⁺+11Fe²⁺

Greenockite (CdS)

Cadmium metal and compounds are mainly used for alloys, coatings,batteries and plastic stabilizers. There are no mines specifically forcadmium extraction. Cadmium sulfide is usually associated with zincsulfides and is recovered as a byproduct of zinc leaching from roastedsulfide concentrates.

Violarite (FeNi₂S₄)

Violarite is a nickel (III) sulfide mineral that is usually associatedwith primary pentlandite nickel sulfide ores.

Reagents

A person skilled in the art will also understand that reagents having athiocarbonyl functional group include, but are not limited to thiourea(Tu), thioacetamide (TA), sodium-dimethyldithiocarbamate (SDDC),ethylene trithiocarbonate (ETC) and thiosemicarbazide (TSCA).

A non-exhaustive list of additional compounds having a thiocarbonylfunctional group is: isothiourea; N—N′ substituted thioureas;2,5-dithiobiurea; dithiobiuret; Thiosemicarbazide purum,Thiosemicarbazide; Methyl chlorothiolformate; Dithiooxamide;Thioacetamide; 2-Methyl-3-thiosemicarbazide;4-Methyl-3-thiosemicarbazide; Vinylene trithiocarbonate purum; Vinylenetrithiocarbonate; 2-Cyanothioacetamide; Ethylene trithiocarbonate;Potassium ethyl xanthogenate; Dimethylthiocarbamoyl chloride;dimethyldithiocarbamate; S,S′-Dimethyl dithiocarbonate; Dimethyltrithiocarbonate; N,N-Dimethylthioformamide;4,4-Dimethyl-3-thiosemicarbazide; 4-Ethyl-3-thiosemicarbazide;O-Isopropylxanthic acid; Ethyl thiooxamate; Ethyl dithioacetate;Pyrazine-2-thiocarboxamide; Diethylthiocarbamoyl chloride;diethyldithiocarbamate; Tetramethylthiuram monosulfide;Tetramethylthiuram disulfide; Pentafluorophenyl chlorothionoformate;4-Fluorophenyl chlorothionoformate; O-Phenyl chlorothionoformate;O-Phenyl chlorothionoformate; Phenyl chlorodithioformate;3,4-Difluorothiobenzamide; 2-Bromothiobenzamide; 3-Bromothiobenzamide;4-Bromothiobenzamide; 4-Chlorothiobenzamide; 4-Fluorothiobenzamide;Thiobenzoic acid; Thiobenzamide; 4-Phenylthiosemicarbazide; O-(p-Tolyl)chlorothionoformate; 4-Bromo-2-methylthiobenzamide;3-Methoxythiobenzamide; 4-Methoxythiobenzamide;4-Methylbenzenethioamide; Thioacetanilide; Salicylaldehydethiosemicarbazone; Indole-3-thiocarboxamide; S-(Thiobenzoyl)thioglycolicacid; 3-(Acetoxy)thiobenzamide; 4-(Acetoxy)thiobenzamide; methylN′-[(e)-(4-chlorophenyl)methylidene]hydrazonothiocarbamate;3-Ethoxythiobenzamide; 4-Ethylbenzene-1-thiocarboxamide; tert-Butyl3-[(methylsulfonyl)oxy]-1-azetanecarboxylate; Diethyldithiocarbamicacid; 2-(Phenylcarbonothioylthio)propanoic acid; 2-Hydroxybenz aldehydeN-ethylthiosemicarbazone;(1R,4R)-1,7,7-Trimethylbicyclo[2.2.1]heptane-2-thione; Tetraethylthiuramdisulfide; Tetraethylthiuram disulfide;4′-Hydroxybiphenyl-4-thiocarboxamide; 4-Biphenylthioamide; Dithizone;4′-Methylbiphenyl-4-thiocarboxamide; tetraisopropylthiuram disulfide;Anthracene-9-thiocarboxamide; Phenanthrene-9-thiocarboxamide; Sodiumdibenzyldithiocarbamate; and 4,4′-Bis(dimethylamino)thiobenzophenone.Such agents are ready available from, for example, Sigma Aldrich.

Each of Tu, TA, SDDC, ETC and TSCA feature a thiocarbonyl functionalgroup having a sulfur that 1) bears a partial negative charge, 2) bearsnegative electrostatic potential surface, and 3) has an emptyπ*-antibonding orbital as its lowest unoccupied molecular orbital(LUMO). Accordingly, the skilled person may reasonably expect that otherreagents, including those additional reagents listed above, that sharesuch criteria and are sufficiently soluble in water may be useful in theperformance of the methods disclosed herein (provided that they do notcomplex with the metal or iron oxidant to from precipitates). It will bewithin the purview of the skilled person to identify potentially usefulreagents and test them to determine efficacy with any particular ore, ifany at all.

For example, Tu has a thiocarbonyl functional group with the sulfurbearing a partial charge of −0.371, a negative electrostatic potentialaround the Sulfur, and π*-antibonding orbital as its LUMO. Hence,thiourea satisfies all three criteria and has demonstrated catalyticeffect.

TA has a similar structure as Tu, but with a CH₃ side chain instead ofNH₂. It has a thiocarbonyl functional group with the sulfur bearing apartial charge of −0.305, which is slightly lower than that for Tu, anegative electrostatic potential around the sulfur, and a π*-antibondingorbital as its LUMO. Accordingly, TA also satisfies all three criteriaand has demonstrated catalytic effect.

ETC differs from Tu and TA as it does not contain any thioamide group.It has a thiocarbonyl functional group with the two sulfur atomsσ-bonded to carbon as the side chain. The sulfur in the thiocarbonylgroup bears a partial charge of −0.122, which is much lower than Tu, anegative electrostatic potential around the Sulfur, and π*-antibondingorbital as its LUMO. Accordingly, ETC also satisfies all three criteriaand has demonstrated catalytic effect.

In comparison, urea has a carbonyl functional group with a C═O bondinstead of C═S. The oxygen in the C═O bond bears a partial charge of−0.634 and a negative electrostatic potential around it, which is verysimilar to the sulfur atom in Tu. However, its LUMO does not containπ*-antibonding. Accordingly, urea is not predicted to have a catalyticeffect in metal leaching, which is confirmed in respect of chalcopyriteby the results of the stirred reactor experiment shown in FIG. 22.

Carbon disulfide (CS₂) contains two thiocarbonyl functional groups.Although the sulfur atoms of each functional group contain aπ*-antibonding orbitals as their LUMO, they bear a partial positivecharge of +0.012. Therefore, CS₂ is not predicted to have catalyticeffect, which is confirmed in respect of chalcopyrite by the results ofthe stirred reactor experiment shown in FIG. 23.

Of course, the reagent should also be water soluble. ETC, for example,is only sparingly soluble in water, which may explain why it appearsless effective than Tu in leaching copper from chalcopyrite.

Preferentially, the reagent will not form complexes/precipitate withFe²⁺/Fe³⁺ ions. TSCA, for example, is able to form a red-color complexwith Fe³⁺ in solution, which may explain why it is less effective thanTu in leaching copper from chalcopyrite.

The reagent also should not complex/precipitate with target metal ionssuch as Cut, Cu²⁺, Cd²⁺, or Ni²⁺. Dithiooxamide forms an insolublecomplex with copper ions and therefore cannot be used for the leachingof copper sulfide minerals, whereas TA complexes with Cd²⁺ ions to forman insoluble complex and therefore cannot be used for leaching cadmiumsulfide minerals such as greenockite.

Again, the skilled person will appreciate that not all compoundscomprising a thiocarbonyl functional group will be useful in increasingthe rate of metal extraction from a metal sulfide. Furthermore, theskilled person will appreciate that a reagent that works to increase therate of extraction of metal from one metal sulfide may not be useful toincrease the rate of extraction of a metal from a different metalsulfide. Again, it will be within the purview of the skilled person toidentify potentially useful reagents and test them to determine efficacywith any particular ore, if any at all.

Formamidine Disulfide (FDS)

Formamidine disulfide (FDS) is generated by oxidation of Tu. In thepresence of an oxidant such as ferric sulfate, Tu will oxidize partiallyto formamidine disulfide (FDS) according to the following half-cellreaction:

2SC(NH₂)₂→[(NH₂)₂CS]2²⁺+2e ⁻

FDS contains no thiocarbonyl functional group but a sulfur-sulfur sigmabond instead. An equilibrium exists between FDS and Tu in a ferricsulfate solution, such that a leach solution prepared with FDS ratherthan Tu will provide the Tu necessary for catalysis of the metal sulfideleach. That is, a molecule of FDS will dissociate into two molecules ofTu upon dissolution in the ferric sulfate leach solution. Accordingly, aleaching solution employing Tu as the reagent having the thiocarbonylfunctional group may be effectively be prepared using either Tu or FDS.

The skilled person will understand that, due to this equilibrium, theconcentration of Tu (and FDS) may fluctuate over time. Accordingly,“concentration” as used herein to refer to the concentration of Tu inthe leach solution relates to the amount of Tu present in the solutionas if all FDS in the solution was dissociated into Tu (i.e ignoringinterconversion between the two forms). Similarly, “concentration” asused herein to refer to the concentration of FDS in the leach solutionrelates to the amount of FDS present in the solution as if all Tu in thesolution was converted into FDS (i.e ignoring interconversion betweenthe two forms).

“Initial concentration” is used herein to refer to the initialconcentration of the reagent at the time the leach solution is appliedto the ore sample. However, the skilled person will understand that theconcentration of the reagent may diminish over time (e.g. throughprecipitation or decay) as the solution percolates through the column orthe heap. Accordingly, the skilled person will appreciate that theprocesses disclosed herein should work to increase the rate of metalextraction from the metal sulfide provided that the concentration of thereagent is within a suitable range during some portion of thepercolation through the ore.

In the presence of FDS and ferric sulphate (or another suitableoxidant), the anodic dissolution of a copper sulfide mineral such aschalcopyrite may proceed according to the following two reactions, withoxidation of the chalcopyrite by either FDS or ferric, respectively:

CuFeS₂(s)+2[(NH₂)₂CS]₂SO₄(aq)→CuSO₄(aq)+FeSO₄(aq)+2S⁰(s)+4SC(NH₂)₂(aq)

CuFeS₂(s)+2Fe₂(SO₄)₃(a)→CuSO₄(a)+5FeSO₄(a)+2S⁰(s)

After chalcopyrite is oxidized, and the copper is leached from theconcentrate, it is desirable to recover the copper from the pregnantleach solution.

The methods disclosed herein involve two basic steps, namely, leachingand metal recovery (e.g. by SX-EW). The leaching process may be carriedout as a percolation leach (such as a heap leach), a vat leach, or atank leach as is known in the field.

For the purposes of this disclosure, the words “containing” and“comprising” are used in a non-limiting sense to mean that itemsfollowing the word are included, but items not specifically mentionedare not excluded. A reference to an element by the indefinite article“a” does not exclude the possibility that more than one of the elementsis present, unless the context clearly requires that there be one andonly one of these elements.

A “percolation leach”, as used herein, refers to the selective removalof a mineral by causing a suitable solvent to seep into and through amass or pile of material containing the desired soluble mineral, e.g. acolumn leach or a heap leach.

A “column leach”, as used herein, refers to leaching through the use ofa long narrow column in which ore sample and solution are in contact formeasuring the effects of typical variables encountered in actual heapleaching.

A “heap leach”, as used herein, is a process through which metals areextracted from the ore in which they are found, i.e. withoutbeneficiation. A heap leach is often chosen for its efficiency andcost-effectiveness. After being removed from the ground, ore istypically sent through a crusher to break the ore down into smallerparticles (although heap ores can be “run-of-mine” in which the ore isleached in an “as-blasted” state with no further crushing). Heap oresmay be the product of primary, secondary, or tertiary crushing.Traditionally, the crushed particles are then “heaped”, or “stacked”into a large pile.

A persistent cause of failure of heap leach operations is the presenceof excess fines in the materials placed on the pad. Excess fines resultsin a low permeability material and thus the seepage rate of thelixiviant is too slow, or ore-solution contact is insufficient, foreconomic pad operations. Accordingly, the efficiency of a heap leach maybe increased by agglomeration after crushing. “Agglomeration”, as usedherein, refers to a technique that binds together material fines orparticles to create a larger product. Agglomeration may be achieved bydifferent methods known in the art. Typically, heap leach agglomerationis performed in a drum agglomerator with sulfuric acid and no binder, oron conveyor belts with acid sprayed onto the ore at drop points.

The heap is irrigated with a solution that is dependent upon the type ofore being extracted. Acid for the leach will preferably be generated bybacteria using processes known in the art. Alternatively, additionalacid could be added as necessary.

The irrigated solution is allowed to percolate through the ore, anddrain to the bottom of the heap. The ore pile sits over an impermeablelayer, such as plastic sheet, which collects the pregnant leach solutionas it drains through and directs it to a collection pond. Once thesolution is collected, it is pumped to a recovery plant to extract thecopper by solvent extraction and electrowinning (SX-EW).

Applying the methods disclosed herein to a heap leach, ore containing anappropriate sulfide mineral is leached selectively in the presence ofthe acid sulfate and the reagent having a thiocarbonyl functional group.The concentration of the reagent having a thiocarbonyl functional groupin the leach solution may be about 30 mM or perhaps even higher. Theskilled person will understand that it is only necessary that thereagent concentration be within a range sufficient to increase the leachrate of the metal sulfide.

Moreover, while the results presented herein indicate that reagentconcentrations of about 30 mM or less are sufficiently low to facilitatethe leaching of metal from a particular metal sulfide, 30 mMconcentrations may not be economically feasible at the present time.Accordingly, it may be preferable to use lower concentrations of reagentthat are feasible from economic and operational points of view, e.g.about 20 mM or less, about 10 mM or less, about 5 mM or less, about 4 mMor less, about 3 mM or less, about 2 mM or less, about 1.5 mM or less,about 1 mM or less, about 0.9 mM or less, about 0.8 mM or less, about0.7 mM or less, about 0.6 mM or less, about 0.5 mM or less, about 0.4 mMor less, 0.3 mM or less, or about 0.2 mM.

Accordingly, the concentration of the reagent in the acidic sulfatesolution may in the range of about 0.2 mM to about 0.3 mM, about 0.2 mMto about 0.4 mM, about 0.2 mM to about 0.5 mM, about 0.2 mM to about 0.6mM, about 0.2 mM to about 0.7 mM, about 0.2 mM to about 0.8 mM, about0.2 mM to about 0.9 mM, about 0.2 mM to about 1.0 mM, about 0.2 to about1.5 mM, about 0.2 to about 2.0 mM, about 0.2 to about 2.5 mM, about 0.2to about 3 mM, about 0.2 to about 4 mM, about 0.2 to about 5 mM, about0.2 to about 10 mM, about 0.2 to about 20 mM, or about 0.2 to about 30mM.

The leaching process may be run at temperatures between 0° C. (i.e. thefreezing point of water) and 80° C. However, the process would typicallybe carried out at ambient temperature and atmospheric pressure.

Following the leaching process, copper can be extracted from the leachsolution. After a solid-liquid separation, i.e. drainage of the pregnantleach solution containing the copper from the heap, the pregnantsolution is preferably subjected to conventional solvent extraction andelectrowinning to produce pure copper cathodes according to thefollowing overall reaction:

SX-EW: CuSO₄(a)+H₂O(l)→Cu(s)+H₂SO₄(a)+½O₂(g)

Reagents having a thiocarbonyl functional group in the pregnant leachsolution should not present any problem in the electrowinning operationand, as a matter of fact, may even be useful as a leveling agent.Raffinate containing thiourea may then be recirculated to the heap forfurther leaching. The recirculated leach solution may also besupplemented with thiourea to arrive at the desired initial thioureaconcentration for the leach.

EXAMPLES

To facilitate the extraction of metal ions from the minerals listedabove, reagents having a thiocarbonyl functional group were added toacidic ferric sulfate solutions as catalysts. In the experimentsdisclosed herein, it was found that the reagents that containthiocarbonyl functional groups have positive catalytic effect on theextraction of the minerals. Among all the reagents, Tu consistentlyprovided the highest catalytic performance. Accordingly, Tu was the mostheavily studied reagent of those identified. However, the results ofexperiments with other reagents having thiocarbonyl functional groupsare provided to compare their catalytic effects. FDS which does notcontain a thiocarbonyl functional group but has comparable catalyticeffect as Tu was studied as a special case due to its equilibrium withTu.

Leaching reactions were carried out at atmospheric pressure on a varietyof ore compositions, reagent concentrations, ferric concentrations, andunder various other conditions, as described below.

Example 1 Extraction of Copper from Chalcopyrite Using Thiourea Example1.1

The effect of thiourea on the electrochemical behavior of a chalcopyriteelectrode was studied in a conventional 3-electrode glass-jacketed cell.A CuFeS₂ electrode was using as working electrode, a saturated calomelelectrode (SCE) was used as reference, and a graphite bar was used ascounter-electrode. The CuFeS₂ electrode was polished using 600 and 1200grit carbide paper. All experiments were conducted at 25° C. using acontrolled temperature water bath. The electrolyte composition was 500mM H₂SO₄, 20 mM Fe₂SO₄ and 0-100 mM thiourea. Before starting anymeasurement, solutions were bubbled with N₂ for 30 minutes to reduce theconcentration of dissolved O₂. Open circuit potential (OCP) was recordeduntil changes of no more than 0.1 mV/min were observed. After a steadyOCP value was observed, electrochemical impedance spectroscopy (EIS) wasconducted at OCP using a 5 mV a.c. sinusoidal perturbation from 10 kHzto 10 mHz. Linear polarization resistance (LPR) tests were alsoconducted using a scan rate of 0.05 mV/s at ±15 mV from OCP.

Linear potential scans were conducted at electrode potentials ±15 mVfrom the OCP measured at each thiourea concentration. All scans showed alinear behavior within the electrode potential range analyzed. Anincrease in the slope of the experimental plots was observed withincreasing thiourea concentration. The slope of these curves was used toestimate the value of the polarization resistance (R_(ct)) at eachconcentration. These values were then used to estimate the values of thedissolution current density using equation 1:

$\begin{matrix}{i_{dissol} \approx \frac{RT}{{nFR}_{ct}}} & {{Eq}.\mspace{11mu} (1)}\end{matrix}$

FIG. 1 shows the effect of thiourea on the dissolution current densityand mixed potential of the CuFeS₂ electrode, and indicates that amaximum dissolution current density was achieved when thioureaconcentration is 30 mM. Increasing thiourea concentration to 100 mMresulted in a decrease in the current density and mixed potential of theCuFeS₂ electrode. Moreover, after immersing the CuFeS₂ electrode in the100 mM thiourea solution, a copper-like film was observed on the surfaceof the electrode, which film could only be removed by polishing theelectrode with carbide paper.

Example 1.2

FIG. 2 is a bar graph showing the effect of initial thiourea or FDSconcentration on the electrochemical dissolution of a chalcopyriteelectrode in sulfuric acid solution at pH 2 and 25° C. A concentrationof 10 mM thiourea in the leach solution resulted in a six fold increasein dissolution rate compared to no thiourea, and a concentration of 5 mMFDS resulted in a six fold increase relative to 10 mM thiourea. Aconcentration of 10 mM thiourea in leach solution also containing 40 mMFe(III) resulted in a thirty fold increase in dissolution rate comparedto 40 mM Fe(III) alone.

Example 1.3

A column leach of different acid-cured copper ores was conducted withthiourea added to the leach solution. A schematic description of thecolumn setup is shown in FIG. 3. The column diameter was 8.84 cm, thecolumn height was 21.6 cm, and the column stack height was 15.9 cm. Theirrigation rate was 0.77 mL/min or 8 L/m²/h. The pregnant leach solutionemitted from these columns was sampled for copper every 2 or 3 daysusing Atomic Absorption Spectroscopy (AAS).

The specific mineralogical composition of these ores are provided inTable 1. The Cu contents of Ore A, Ore B, and Ore C were 0.52%, 1.03%,and 1.22% w/w, respectively. Prior to leaching, ore was “acid cured” toneutralize the acid-consuming material present in the ore. That is, theore was mixed with a concentrated sulfuric acid solution composed of 80%concentrated sulfuric acid and 20% de-ionized water and allowed to sitfor 72 hours. For one treatment using Ore C, thiourea was added to thesulfuric acid curing solutions.

The initial composition of the leaching solutions included 2.2 g/L Fe(i.e. 40 mM, provided as ferric sulfate) and pH 2 for the controlexperiment, with or without 0.76 g/L thiourea (i.e. 10 mM). The initialload of mineral in each column was 1.6 to 1.8 kg of ore. The superficialvelocity of solution through the ore column was 7.4 L m⁻² h⁻¹. The pHwas adjusted using diluted sulfuric acid. These two columns weremaintained in an open-loop or open cycle configuration (i.e. no solutionrecycle) for the entire leaching period.

The results of leaching tests on the Ore A, Ore B and Ore C are shown inFIGS. 4, 5, and 6, respectively. The presence of thiourea in thelixiviant clearly has a positive effect on the leaching of copper fromthe chalcopyrite. On average, the leaching rate in the presence ofthiourea was increased by a factor of 1.5 to 2.4 compared to the controltests in which the leach solutions did not contain thiourea. As of thelast time points depicted in FIGS. 4 to 6, copper extractions forcolumns containing Ore A, Ore B, and Ore C leached with a solutioncontaining sulfuric acid and ferric sulfate alone, without addedthiourea, were 21.2% (after 198 days), 12.4% (after 50 days), and 40.6%(after 322 days), respectively. With 10 mM of added thiourea, theseextractions were 37.9%, 32.0%, and 72.3%, respectively.

Referring to FIG. 6, 2 mM Tu was added to the leach solution originallycontaining no Tu from day 322 onward, after which the leach rateincreased sharply. From day 332 to day 448, the copper leached from thiscolumn increased from 40% to 58%, and rapid leaching was maintainedthroughout that period.

The averages for the last 7 days reported in FIG. 7 indicate that theleaching rate for acid-cured Ore C leached in the presence of 10 mMthiourea is 3.3 higher than for acid-cured Ore C leached in the absenceof thiourea, and 4.0 times higher than acid-cured and thiourea-cured OreC leached in the absence of thioruea.

FIG. 8 shows the effect of Tu on solution potential. All potentials arereported against a Ag/AgCl (saturated) reference electrode. The solutionpotential of the leach solutions containing thiourea was generallybetween 75 and 100 mV lower than the solution potential of leachsolution that did not include thiourea. Lower solution potentials areconsistent with thiourea working to prevent the passivation ofchalcopyrite.

Example 1.4 Bottle Roll Leaching

“Bottle roll” leaching experiments in the presence of variousconcentrations of thiourea were conducted for coarse Ore A and Ore B.The tests were conducted using coarsely crushed (100% passing ½ inch)ore.

Prior to leaching, the ore was cured using a procedure similar to whatwas performed on the ore used in the column leaching experiments. Theore was mixed with a concentrated sulfuric acid solution composed of 80%concentrated sulfuric acid and 20% de-ionized water and allowed tosettle for 72 hours to neutralize the acid-consuming material present inthe ore. For several experiments, different concentrations of thioureawere added to the ore using the sulfuric acid curing solutions.

The bottles used for the experiments were 20 cm long and 12.5 cm indiameter. Each bottle was loaded with 180 g of cured ore and 420 g ofleaching solution, filling up to around one third of the bottle'svolume.

The leaching solution from each bottle was sampled at 2, 4, 6 and 8hours, and then every 24 hours thereafter. Samples were analyzed usingatomic absorption spectroscopy (AAS) for their copper content.

The conditions for the bottle roll experiments are listed in Table 2.Experiments #1 to #6 were conducted using only the original addition ofthiourea into the bottles. For experiments #7 to #11, thiourea was addedevery 24 hours to re-establish the thiourea concentration.

A positive effect of thiourea on copper leaching was observed. For thecoarse ore experiments, a plateau was not observed until after 80 to 120hours. Thiourea was added periodically to the coarse ore experiments,yielding positive results on copper dissolution.

The effect of different concentrations of thiourea in the leach solutionon the leaching of coarse ore (experiments #1 to #11 as described inTable 2) is shown in FIGS. 9 and 10.

For ore B, thiourea was periodically added every 24 hours tore-establish the thioruea concentration in the system and thus betteremulate the conditions in the column leach experiments. As may beobserved from FIG. 9, 8 mM and 10 mM thiourea yielded higher copperdissolution results than the other thiourea concentrations tested forore A. A plateau in dissolution is not observed until afterapproximately 120 hours, which varied with thiourea concentration asshown in FIG. 9.

TABLE 1 Mineral Ideal Formula Ore A Ore B Ore C Actinolite Ca₂(Mg,Fe²⁺)₅Si₈O₂₂(OH)₂ — 1.8 — Biotite K(Mg, Fe²⁺)₃AlSi₃O₁₀(OH)₂ — 4.2 —Calcite CaCO₃ — 19.3  — Chalco- CuFeS₂ 1.4 3.5 2.6 pyrite Clino- (Mg,Fe²⁺)₅Al(Si₃Al)O₁₀(OH)₈ — 15.0  — chlore Diopside CaMgSi₂O₆ — 3.5 —Galena PbS — — 0.1 Gypsum CaSO₄2H₂O — 1.2 — Hematite α-Fe₂O₃ — 0.2 —K-feldspar KAlSi₃O₈ 17.9 10.8  — Kaolinite Al₂Si₂O₅(OH)₄ 2.3 — 2.3Magnetite Fe₃O₄ — 0.8 — Molyb- MoS₂ <0.1 — — denite MuscoviteKAl₂AlSi₃O₁₀(OH)₂ 21.9 6.0 41.6  Plagio- NaAlSi₃O₈—CaAlSi₂O₈ 13.6 25.4 — clase Pyrite FeS₂ 2.3 — 8.0 Quartz SiO₂ 40.0 8.3 44.4  Rutile TiO₂ 0.5— 0.9 Siderite Fe²⁺CO₃ — 0.1 — Total 100 100    100   

As may be observed from FIG. 9, 5 mM thiourea yielded higher copperdissolution results than the other thiourea concentrations tested forore B. As with ore A, a plateau in dissolution is not observed untilafter approximately 80 to 120 hours, which varied with thioureaconcentration as shown in FIG. 9. Periodic addition of thiourea resultedin increased copper dissolutions and produced a delay in the dissolutionplateau.

Interestingly, solutions containing 100 mM thiourea did not appear to bemuch more effective on copper extraction than those containing nothiourea, and even worse at some time points. This is consistent withthe results of Deschênes and Ghali, which reported that solutionscontaining ˜200 mM thiourea (i.e. 15 g/L) did not improve copperextraction from chalcopyrite. Thiourea is less stable at highconcentrations and decomposes. Accordingly, it is possible that, wheninitial thiourea concentrations are somewhat higher than 30 mM,sufficient elemental sulfur may be produced by decomposition of thioureato form a film on the chalcopyrite mineral and thereby assist in itspassivation. It is also possible that, at high Tu dosages, some copperprecipitates from solution (e.g. see FIG. 15) to account for some of thelow extraction results.

Example 2 Extraction from Chalcopyrite, Covellite, Chalcocite, Bornite,Enargite, Pentlandite, Violarite, and Greenockite Using Thiourea

The catalytic effect of Tu was further demonstrated in stirred reactortests. All reactors contained 1.9 L of ferric sulfate solution at pH 1.8and total iron concentration of 40 mM. 1 g of mineral samples was usedin each reactor test. These experimental conditions were designed tomaintain an unlimited supply of oxidant.

In order to demonstrate the catalytic effect on chalcopyrite, 100% puresynthetic chalcopyrite was used instead of chalcopyrite concentratewhich contains various impurities. The chalcopyrite was synthesized viaa hydrothermal approach. CuCl, FeCl₃ and thiourea were first mixed witha molar ratio of 1:1:2 and dissolved in 150 mL DI water. The solutionwas transferred to a Teflon-lined reaction vessel and heated up to 240°C. for 24 hours. At the end of the reaction, the precipitated powder waswashed with acidic water (pH=1) and dried in air at room temperature.XRD analysis in showed that the synthetic chalcopyrite was free of anyimpurities compared with chalcopyrite mineral concentrate. Thissynthetic chalcopyrite was used in all the tests carried out in stirredreactors as disclosed herein.

The covellite mineral used in the experiment disclosed herein was alsosynthesized via a hydrothermal approach. CuCl and Tu were mixed with amolar ratio of 1:1 and dissolved in 150 mL DI water. The solution wastransferred to a Teflon-lined reaction vessel and heated up to 220° C.for 24 hours. The synthesized CuS was acid-washed and dried in air. XRDanalysis showed that it had 100% purity with no interference of otherspecies.

The chalcocite mineral sample used in the experiments disclosed hereinwas 100% pure natural mineral.

The bornite mineral used in the experiments disclosed herein wasobtained from Butte, Mont. with copper content of 58.9% based onICP-AES. XRD analysis showed that the mineral contains 76.8% bornite,8.1% chalcopyrite, 6.3% pyrite, 5.8% tennatite and 3.0% enargite. Thecopper content calculated from XRD was 55.6%, which is relativelyconsistent with the chemical assay.

The enargite mineral used in the experiments disclosed herein was in theform of an enargite concentrate, which contained approximately 70%enargite (34% copper) according to XRD analysis.

The Greenockite mineral used in this experiment was synthesized via ahydrothermal approach. CdCl₂ and thiourea were mixed with a molar ratioof 1:1 and dissolved in 100 mL DI water. The solution was transferred toa Teflon-lined reaction vessel and heated up to 150° C. for 24 hours.The synthesized CdS was acid-washed and dried in air. XRD analysisshowed that it has 100% purity with no interference of other species.

TABLE 2 List of bottle roll leaching experiments involving Ore A and OreB. Experiment Brief description of experimental conditions #1 Coarse oreA, 0 mM Tu in solution, 40 mM ferric in solution, acid curing, no Tureplenishment #2 Coarse ore A, 2 mM Tu in solution, 40 mM ferric insolution, acid curing, no Tu replenishment #3 Coarse ore A, 4 mM Tu insolution, 40 mM ferric in solution, acid curing, no Tu replenishment #4Coarse ore A, 6 mM Tu in solution, 40 mM ferric in solution, acidcuring, no Tu replenishment #5 Coarse ore A, 8 mM Tu in solution, 40 mMferric in solution, acid curing, no Tu replenishment #6 Coarse ore A, 10mM Tu in solution, 40 mM ferric in solution, acid curing, no Tureplenishment #7 Coarse ore B, 0 mM Tu in solution, 40 mM ferric insolution, acid curing #8 Coarse ore B, 1 mM Tu in solution, 40 mM ferricin solution, acid curing, periodic addition of Tu to replenish 1 mMconcentration in solution #9 Coarse ore B, 5 mM Tu in solution, 40 mMferric in solution, acid curing, periodic addition of Tu to replenish 5mM concentration in solution #10  Coarse ore B, 10 mM Tu in solution, 40mM ferric in solution, acid curing, periodic addition of Tu to replenish10 mM concentration in solution #11  Coarse ore B, 100 mM Tu insolution, 40 mM ferric in solution, acid curing, periodic addition of Tuto replenish 100 mM concentration in solution

The violarite used in the experiments disclosed herein was naturalviolarite mineral that contains 15.8% Ni according to ICP-AES. XRDanalysis showed that the mineral had approximately 42% violarite and13.1% NiSO₄.6H₂O.

The sulfur on thiocarbonyl groups contains a lone electron pair and afilled π-orbital which can be used for donor-acceptor type bonding witha transition metal, together with a π*-antibonding orbital that couldpotentially accept the back-donation of electrons from the filledd-orbitals on the transition metal. Accordingly, without wanting to bebound by theory, it is suspected that the interaction between thesurface ion and the thiocarbonyl functional group, especially backdonation from metal to ligand, is responsible for the catalytic effect.Moreover, it is suspected that the catalytic effect should be morepronounced for the transition metals with higher d-electron numbers,with the catalytic effect being most pronounced for minerals with d¹⁰electronic configuration

FIG. 11 shows that Tu catalyzes the leaching of common copper sulfideminerals, including chalcopyrite, covellite, chalcocite, and bornite,which all contain Cu(I). After 96 hours of leaching, chalcopyriteextraction reaches 64.1% with 2 mM of Tu compared to 21.1% without Tu;covellite extraction reaches 74.4% with 2 mM of Tu compared to 7.2%without Tu; chalcocite extraction reaches 85.6% with 2 mM of Tu comparedto 65.1% without Tu; bornite extraction reaches 91.4% with 2 mM of Tucompared to 56.7% without Tu.

Like Cu(I), Cd(II) also contains the d¹⁰ electronic configuration. FIG.12 shows that leaching of CdS mineral is significantly enhanced with theaddition of Tu. With Tu, the extraction of cadmium reaches 100% at 48hours whereas extraction in the noncatalyzed reaction plateaued at 47%after 96 hours.

The copper ion in the enargite mineral has fewer d-electrons than otherprimary and secondary sulfides, and thus it may be expected that thecatalytic effect should be slower than that for Cu(I) minerals.Nevertheless, the results shown in FIG. 13 clearly demonstrate thatleaching with a leach solution comprising an initial concentration of 2mM Tu increases the leach rate of copper from enargite compared to acontrol without Tu, which did not show any significant extraction after96 hours of leaching.

Minerals that contain transition metal ions with d⁷ electronicconfiguration, such as Ni(III), may also undergo catalyzed leaching withthe addition of Tu. Similar to Cu(II), as Ni(III) is the highest stableoxidation state with 7 d-electrons, the catalytic effect is not expectedto be as dramatic as for d¹⁰ minerals. Referring to FIG. 14, leachingwith a leach solution comprising an initial concentration of 2 mM Tuincreases the leach rate of nickel from violarite compared to a controlwithout Tu.

Results of leaching experiments referred to in Example 2 are summarizedin Table 3, in which the extraction percentages under non-catalyzed andcatalyzed conditions (with an initial Tu concentration of 2 mM) arecompared.

TABLE 3 Comparisons of reactor leaching for various minerals underuncatalyzed and 2 mM Tu catalyzed conditions 96-Hour Extraction 96-HourExtraction Mineral (No thiourea) (2 mM thiourea) Chalcopyrite, CuFeS₂21.1% 64.1% Covellite, CuS 6.8% 74.4% Chalcocite, Cu₂S 65.1% 85.5%Bornite, Cu₅FeS₄ 56.7% 91.4% Greenokite, CdS 46.5% 100.0% Enargite,Cu₃AsS₄ 2.1% 10.0% Violarite, FeNi₂S₄ 13.0% 22.2%

Example 3 Reagent Dosage

Optimum dosage of reagent may increase the efficiency of leaching.First, at certain concentrations, the reagent may form an insolublecomplex with the metal ion of interest and precipitate. For example, Tucan form an insoluble complex with Cu(I) ions at a 3:1 molar ratio. Aprecipitation test was performed to examine the concentration range atwhich Cu-Tu complex precipitation may occur. 20 mL of Cu solution wasdivided into several identical portions followed by the addition ofvarious Tu dosage (i.e. 0 to 60 mM). The solution was stirred for 24hours, and the Cu remaining in the solution phase was analyzed by AAS.The results are shown in FIG. 15, plotted as the percentage of Curemaining.

Second, heap leaching of metal sulfides is based on a bioleachingmechanism, an excessive amount of reagent may be detrimental tobioleaching microbes. For example, bacteria commonly used forbioleaching, such as Acidithiobacillus ferrooxidans andAcidithiobacillus thiooxidans, have very slow growth in a solutioncontaining 10 mM Tu, and cannot survive at 100 mM Tu.

Third, with respect to Tu specifically, ferric reacts with Tu andconverts it to FDS (see Hydrometallurgy 28, 381-397 (1992)). Althoughthe reaction is reversible under certain conditions, a highconcentration of FDS tends to decompose irreversibly into cyanamide andelemental sulfur (see J Chromatogr 368, 444-449).

2Tu+2Fe³⁺↔FDS+2Fe²⁺+2H⁺

FDS→Tu+cyanimide+S

Therefore, over-addition of Tu in the lixiviant may cause the loss ofFe³⁺ and Tu due to oxidation and decomposition. The irreversibledecomposition of FDS has been observed when adding 4 mM of Tu into a 40mM ferric sulfate solution at pH 1.8.

To further investigate the effect of Tu dosage on copper extraction,stirred reactor tests were performed using 1 g of synthetic chalcopyritein 1.9 L of 40 mM ferric sulfate solution at pH 1.8 with various initialTu concentrations. The treatments were run for 172 hours to approachmaximum extraction. The results are presented in FIG. 16, and showsthat, for 1 g of chalcopyrite, higher Tu dosage results in fasterleaching kinetics among the Tu concentrations tested.

For Tu dosages of 5 mM and under, the initial 40 mM ferric sulfatesolution can be considered as a sufficient supply of oxidant. However,for higher dosages such as 10 mM and 20 mM of Tu, extra ferric (in 1:1ratio with Tu) had to be added to the solution to allow the oxidation ofTu to FDS. For 10 mM Tu, an extra 10 mM of Fe³⁺ was added at time zero.For 20 mM Tu, an extra 20 mM of Fe³⁺ was added at 72 hours, which led tothe continuation of extraction as shown in FIG. 16.

The Tu dosage vs. Cu extraction at 172 hours is plotted in FIG. 17. Aninitial Tu dosage up to 5 mM appears to have the most pronounced effecton the dissolution of Cu.

As indicated above, in previous shakeflask tests with acidic solutions(pH 1.8) containing various concentrations of Fe³⁺ and Cu²⁺ ions, slightprecipitation occurred upon the addition of 4 mM of Tu due to thedecomposition of FDS. Accordingly, concentrations of Tu concentrationbelow 4 mM may avoid such precipitation. A series of shakeflask testswere performed on solutions containing initial concentrations of 2 mM Tuand various concentrations in a matrix containing Fe³⁺ (0-100 mM) andCu²⁺ (0-50 mM) in order to identify concentration ranges of [Fe³⁺ ] and[Cu²⁺] that do not result in Cu complex precipitation. The resultsshowed that no precipitation and no loss of Cu from the solution phaseresulted using 2 mM of Tu in this wide range of Fe and Cu matrixconcentrations.

Example 4 Alternative Reagents

The catalytic effect of several other reagents having a thiocarbonylfunctional group was examined on the leaching of synthetic chalcopyrite,covellite, bornite, and enargite.

Experiments were carried out in stirred reactors containing 40 mM ferricsulfate solution at pH 1.8. 1 g of chalcopyrite or covellite was addedto the reactors along with an initial concentration of 2 mM of variousthiocarbonyl reagents including Tu, TA, SDDC, ETC and TSCA. The Cuextraction curves for chalcopyrite, covellite, bornite, and enargiteusing all or a subset of the above reagents are shown in FIGS. 18, 19,20, and 21.

From FIGS. 18 to 21, it is clear that each of these further reagentsthat have a thiocarbonyl functional group show a beneficial effect inthe ferric sulfate leaching of each of chalcopyrite, covellite, borniteand enargite.

FIG. 22 summarizes the results of further stirred reactor tests onchalcopyrite that additionally investigate urea and carbon disulfide.These results confirm that, as expected, neither urea nor carbondisulfide are effective reagents.

Example 5 FDS

The catalytic effect of leaching solutions prepared with FDS onchalcopyrite, bornite, covellite, and chalcocite leaching was determinedin stirred reactor tests. All reactors contained 1.9 L of ferric sulfatesolution at pH 1.8 and total iron concentration of 40 mM. 1 g of mineralsamples was used in each reactor test. An initial FDS concentration of 1mM or an initial Tu concentration of 2 mM Tu was used.

The results from stirred reactor tests shown in FIG. 23, panel a, andFIG. 23, panel b. demonstrate that FDS has comparable efficiency to Tuin the leaching of each of chalcopyrite, bornite, covellite, andchalcocite after 96 hours.

Example 6 Stepwise Closed Loop Bioleaching with Tu

A closed loop bioleach with Tu was conducted. 7 kg of ore containapproximately 0.25% Cu content, mainly in the form of CuFeS₂ was leachedat a flow rate of 1 L/day at an aeration rate of approximately 300mL/min.

The ore was pre-treated with sulfuric acid to leach oxides (e.g.chalcanthite and basic copper salts) using sulfuric acid. After the acidleaching period finished, residual solutions were collected and replacedby a ferrous sulfate solution with nutrients (40 mM FeSO₄, 0.4 g/Lmagnesium sulfate heptahydrate and 0.04 g/L potassium dihydrogenphosphate, with pH adjusted to 1.6-1.8). The ferrous and nutrientssolution was flushed through the column to establish a good habitat forbacterial growth. Inoculation of bacteria showed an increase in the ORPfrom 274 mV to 550 mV within 48 hours. The solution used in this stepand future steps was kept circulating through the column, forming aself-sustaining closed-loop system.

At this stage, the remaining copper source is mainly CuFeS₂. After thebacteria had survived in the column, Tu was progressively added to theleaching solution. As discussed above Tu is converted to FDS at a molarratio of 2:1 in the presence of 40 mM Fe³⁺. Operating potential (ORP)was used as the indicator for bacterial activity, and HPLC was used tomonitor FDS content. From day 0 to day 50, the leaching solutionincluded 40 mM Fe³⁺ with inoculated bacteria (with no Tu addition). Fromday 90 to day 98, a total of 1.878 g of Tu was progressively added, uponwhich the HPLC analysis on the effluent showed that the FDS was beingmaintained at approximately 1.5 mM, and no more Tu was added.

As shown in FIG. 24, the ORP of the effluent was always equal to orhigher than the influent, indicating that bacteria were activelyoxidizing Fe²⁺ to Fe³⁺. The FDS contents were analyzed by HPLC, showingthat approximately 1.5 mM of FDS (equivalent to 3 mM of Tu added)existing in the solution phase without any precipitation being observed.Therefore, it appears that 1.5 mM FDS (3 mM Tu equivalent) may be usedin the solution without precipitation of ferric.

The results of closed loop leaching test are shown in FIG. 25. From day0 to day 50, bacteria were able to maintain high activity and oxidizeFe²⁺ to Fe³⁺. However, with the constant flow rate (1 L/day), theleaching rate was only 1.97 mg Cu/day for the first 50 days. Addition ofTu starting on day 90 increased the Cu extraction rate to 6.54 mg/day,which remained constant after day 98. This indicates that the reagentdid not undergo decomposition and remained effective in the closed-loopsystem.

While specific embodiments of the invention have been described andillustrated, such embodiments should be considered illustrative of theinvention only and not as limiting the invention as construed inaccordance with the accompanying claims.

1.-59. (canceled) 60.-114. (canceled)
 115. A method of recovering atleast one base metal ion from a material containing at least one basemetal sulfide, the method comprising: contacting the material with anacidic solution comprising at least one oxidizing agent and a reagenthaving a thiocarbonyl functional group to extract at least one basemetal ion from the at least one base metal sulfide, wherein theconcentration of the reagent is in the range of 0.2 mM to 100 mM, toproduce a pregnant solution containing the at least one base metal ion;and recovering at least one base metal ion from the pregnant solution.116. The method of claim 115, wherein the concentration of the reagentin the acidic solution is sufficient to increase the rate of the atleast one base metal ion extraction relative to an acidic solution thatdoes not contain the reagent.
 117. The method of claim 115, wherein theat least one oxidizing agent comprises a source of ferric ions.
 118. Themethod of claim 117, wherein the source of ferric ions is Fe³±.
 119. Themethod of claim 117, wherein the source of ferric ions is ferricsulfate.
 120. The method of claim 117, wherein the source of ferric ionscomprises ferric ions generated at least in part by bacteria.
 121. Themethod of claim 115, wherein the acidic solution comprises a ferricsulfate solution.
 122. The method of claim 115, wherein the acidicsolution comprises a ferric solution.
 123. The method of claim 115,wherein the acidic solution comprises a ferric media.
 124. The method ofclaim 115, wherein the acidic solution is replaced with a ferroussulfate solution.
 125. The method of claim 124, wherein the ferroussulfate solution provides a source of Fe²⁺ ions that are activelyoxidized by to Fe³⁺ by bacteria.
 126. The method of claim 115, whereinthe thiocarbonyl functional group of the reagent has a sulfur that bearsa partial negative charge, bears negative electrostatic potentialsurface, and has an empty π*-antibonding orbital as its lowestunoccupied molecular orbital.
 127. The method of claim 115, wherein thereagent is N—N′ substituted thioureas; 2,5-Dithiobimea; Dithiobiuret;Thiosemicarbazide purum; Thiosemicarbazide; Thioacetamide;2-Methyl-3-thiosemicarbazide; 4-Methyl-3-thiosemicarbazide; Vinylenetrithiocarbonate purum; Vinylene trithiocarbonate; 2-Cyanothioacetamide;Ethylene trithiocarbonate; Potassium ethyl xanthogenate;Dimethylthiocarbamoyl chloride; Dimethyldithiocarbamate; Dimethyltrithiocarbonate; N,N-Dimethylthioformamide;4,4-Dimethyl-3-thiosemicarbazide; 4-Ethyl-3-thiosemicarbazide;O-Isopropylxanthic acid; Ethyl thiooxamate; Ethyl dithioacetate;Pyrazine-2-thiocarboxamide; Diethylthiocarbamoyl chloride;Diethyldithiocarbamate; Tetramethylthiuram monosulfide;Tetramethylthiuram disulfide; Pentafluorophenyl chlorothionoformate;4-Fluorophenyl chlorothionoformate; O-Phenyl chlorothionoformate; Phenylchlorodithioformate; 3,4-Difluorothiobenzamide; 2-Bromothiobenzamide;3-Bromothiobenzamide; 4-Bromothiobenzamide; 4-Chlorothiobenzamide;4-Fluorothiobenzamide; Thiobenzoic acid; Thiobenzamide;4-Phenylthiosemicarbazide; O-(p-Tolyl) chlorothionoformate;4-Bromo-2-methylthiobenzamide; 3-Methoxythiobenzamide;4-Methoxythiobenzamide; 4-Methylbenzenethioamide; Thioacetanilide;Salicylaldehyde thiosemicarbazone; Indole-3-thiocarboxamide;S-(Thiobenzoyl)thioglycolic acid; 3-(Acetoxy)thiobenzamide;4-(Acetoxy)thiobenzamide; MethylN′-[(e)-(4-chlorophenyl)methylidene]hydrazonothiocarbamate;3-Ethoxythiobenzamide; 4-Ethylbenzene-1-thiocarboxamide; Tert-butyl3-[(methylsulfonyl)oxy]-1-azetanecarboxylate; Diethyldithiocarbamicacid; 2-(Phenylcarbonothioylthio)propanoic acid; 2-Hydroxybenzaldehyde;N-ethylthiosemicarbazone;(1R,4R)-1,7,7-Trimethylbicyclo[2.2.1]heptane-2-thione; Tetraethylthiuramdisulfide; 4′-Hydroxybiphenyl-4-thiocarboxamide; 4-Biphenylthioamide;Dithizone; 4′-Methylbiphenyl-4-thiocarboxamide; Tetraisopropylthiuramdisulfide; Anthracene-9-thiocarboxamide; Phenanthrene-9-thiocarboxamide;Sodium dibenzyldithiocarbamate; 4,4′-Bis(dimethylamino)thiobenzophenone;or any combination thereof.
 128. The method of claim 115, wherein thematerial is an ore.
 129. The method of claim 115, wherein the materialis a concentrate of the at least one base metal sulfide.
 130. The methodof claim 115, wherein the concentration of the reagent in the acidicsolution is in the range of 0.2 mM to 60 mM.
 131. The method of claim115, wherein the concentration of the reagent in the acidic solution isin the range of 0.2 mM to 50 mM.
 132. The method of claim 115, whereinthe concentration of the reagent in the acidic solution is in the rangeof 0.2 mM to 30 mM.
 133. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 20 mM.
 134. The method of claim 115, wherein the concentrationof the reagent in the acidic solution is in the range of 0.2 mM to 10mM.
 135. The method of claim 115, wherein the concentration of thereagent in the acidic solution is in the range of 0.2 mM to 5 mM. 136.The method of claim 115, wherein the concentration of the reagent in theacidic solution is in the range of 0.2 mM to 4 mM.
 137. The method ofclaim 115, wherein the concentration of the reagent in the acidicsolution is in the range of 0.2 mM to 3 mM.
 138. The method of claim115, wherein the concentration of the reagent in the acidic solution isin the range of 0.2 mM to 2 mM.
 139. The method of claim 115, whereinthe concentration of the reagent in the acidic solution is in the rangeof 0.2 mM to 1.5 mM.
 140. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 1.0 mM.
 141. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.9 mM.
 142. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.8 mM.
 143. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.7 mM.
 144. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.6 mM.
 145. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.5 mM.
 146. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.4 mM.
 147. The method of claim 115, wherein theconcentration of the reagent in the acidic solution is in the range of0.2 mM to 0.3 mM.
 148. The method of claim 115, wherein the at least onebase metal comprises copper and the at least one base metal sulfidecomprises chalcopyrite.
 149. The method of claim 115, wherein the atleast one base metal comprises copper and the at least one base metalsulfide comprises covellite.
 150. The method of claim 115, wherein theat least one base metal comprises copper and the at least one base metalsulfide comprises a copper sulfide of the formula CuxSy wherein the x:yratio is between 1 and
 2. 151. The method of claim 150, wherein thecopper sulfide comprises chalcocite.
 152. The method of claim 150,wherein the copper sulfide comprises djurleite.
 153. The method of claim150, wherein the copper sulfide comprises digenite.
 154. The method ofclaim 115, wherein the at least one base metal comprises copper and theat least one base metal sulfide comprises bornite.
 155. The method ofclaim 115, wherein the at least one base metal comprises copper and theat least one base metal sulfide comprises enargite.
 156. The method ofclaim 115, wherein the at least one base metal comprises cadmium and theat least one base metal sulfide comprises greenockite.
 157. The methodof claim 115, wherein the at least one base metal comprises nickel andthe at least one base metal sulfide comprises pentlandite.
 158. Themethod of claim 115, wherein the at least one base metal comprisesnickel and the at least one base metal sulfide comprises violarite. 159.The method of claim 115, wherein the reagent is thiourea (Tu).
 160. Themethod of claim 115, wherein the reagent is thioacetamide (TA).
 161. Themethod of claim 115, wherein the reagent is sodiumdimethyldithiocarbamate (SDDC).
 162. The method of claim 115, whereinthe reagent is ethylene trithiocarbonate (ETC).
 163. The method of claim115, wherein the reagent is thiosemicarbazide (TSCA).
 164. The method ofclaim 115, wherein the method is a leach.
 165. The method of claim 164,wherein the leach is a percolation leach.
 166. The method of claim 164,wherein the leach is a heap leach.
 167. The method of claim 164, whereinthe leach is a vat leach.
 168. The method of claim 164, wherein theleach is a tank leach.
 169. The method of claim 164 wherein the leach isa column leach.
 170. The method of claim 164, wherein the leach is runat a temperature between 0° C. and 80° C.
 171. The method of claim 115,wherein recovering the at least one metal from the pregnant solutioncomprises solvent extraction and electrowinning.
 172. The method ofclaim 115, further comprising maintaining the operating potential of theacidic solution above 500 mV vs Ag/AgCl.
 173. The method of claim 115,wherein the reagent does not complex/precipitate with the base metalions.